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By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By Cyril Smith

A STUDY of the precipitation-hardening of copper steels1 led the authors to investigate malleable iron containing copper, for the low-carbon ferritic matrix in malleable iron should lend itself admirably to precipitation-hardening. It was soon found that in addition to rendering malleable iron susceptible to improvement by precipitation-hardening treatments, copper has another potentially useful effect as an accelerator of graphitization. Brief mention of these effects was given in a general paper by Lorig and one of the present authors2; the present paper gives further details on the graphitization and precipitation-hardening effects only. After the completion of this work a paper by L. Lykkin3 appeared, confirming the authors' conclusions regarding graphitization.

Jan 1, 1935

By H. L. Hopkins, OHIO, L. W. Kemph, CLEVEL AND, E. V. Ivanso

A balanced system of internal stress is set up in any metallic structure by plastic deformation below the annealing temperature. The internaI stress induced by cold rolling or other fabricating processes is generally recognized and its magnitude is known to be, within certain limits, a function of the degree of cold deformation. The plastic deformation that may accompany rapid cooling from elevated temperatures may not materially change the form of the product. The internal stresses so induced are not, therefore, so evident and nonrecognition of their presence has at times led to engineering failures. The heat treatment of aluminum alloys is a relatively new art and little information is available regarding the magnitude of stresses induced during the commercial heat treatment of aluminum alloy products. As the industrial use of aluminum increases, it is inevitable that products of increasingly large volume will be required. It has become necessary in the present commercial utilization of aluminum to heat-treat castings and forgings weighing hundreds of pounds. One of the factors governing the magnitude of internal stresses arising from rapid cooling from elevated temperatures is that of the size of the structure being treated. Some of the other factors are: cooling rate, elastic properties of the alloy, thermal expansivity, thermal conductivity, and the presence or absence of phase changes. Practically all aluminum alloys have a high coefficient of thermal expansion and some are capable of developing relatively high elastic limits. On the other hand, all commercial aluminum alloys have relatively high thermal conductivities, no invariant phase changes, and the quenching temperatures are comparatively low. In the absence of invariant phase changes involving change in the type of crystallographic lattice, the fundamental basis for the development of internal stress

Jan 1, 1934

By R. S. Sandhu, L. E. Baker, W. Y. Shieh

The finite element method of analysis was introduced by Clough.1 Within a short period, it has developed into a powerful solution technique in structural and continuum mechanics. It has been used to solve boundary value problems in two and three-dimensional elasticity.2-5 Anisotropy and heterogeneity of materials were allowed for at the earliest stages. 1,4 Its application has been extended to allow for cracks,5 creep and viscoelasticity, 6,7 incremental construction,6-8 nonlinear elasticity3,9 effect of joints in rock,10 and to materials with different behavior in tension and compression.7 Because of its versatility, the finite element method has found increasing application in the field of rock mechanics. In fact, several of the more significant developments have been motivated by problems in this field. Elasto-plastic material behavior based on the incremental or rate-type theory of plasticity rather than the deformation theory was considered by Reyes and Deere11 A material law based on a generalization 12 of the Mohr-Coulomb criterion was used. In this chapter, the formulation of material laws developed by Reyes and Deere 11 is used. An efficient analysis procedure, incorporating recent developments in finite element analysis, is developed for plane-strain conditions. To demonstrate its effectiveness, the procedure is applied to three cases, one for which an exact solution is available, another for which field observations had been made, and a third one for which a parallel study was made, using an accepted limit-equilibrium method.

Jan 1, 1970

By Hannu Penttila, Harrison R. Cooper, Olli Mattila

Two approaches to automated operation in metallurgical processes involve the conventional analog control or, more recently, digital control. The expanding use of digital controls applied to process automation is a comparatively new development. Several current technical reports describe digital control systems in concentrator grinding control, 1, 2, flotation, 3 and smelter operations. 4,5 The principle factors influencing trends to digital control are:

Jan 10, 1976

By Joseph M. Dawson

Although there was a very great increase in drilling activity along the fault line of south Texas during 1937 as compared with the previous three or four years, and although seven new fields were discovered, it cannot be said that any major development of importance took place. There were slightly over 550 completions during 1937, of which nearly 350 were oil wells as compared with some 275 the preceding year, of which 169 were oil wells. This increase in drilling activity was occasioned almost entirely by the greater interest in and development of Austin chalk production, principally in Caldwell, Guadalupe and Bastrop Counties. Seven fields were found, of which a brief description is given in Table 3. None has as yet assumed anything of major importance, and it is doubtful whether any will be classified as major discoveries through further development. Two were producing from the Edwards limestone, one from serpentine and the other four from the Austin chalk; the latter may be classified as extremely erratic in nature. The major discouraging feature of the year was the rise and subsequent fall of interest and importance of Austin chalk production. Many new operators came into this district to drill and produce Austin chalk oil toward the first of the year, but in the latter months the production decline from this formation was so rapid and the production so erratic that drilling operations fell off materially and interest largely abated. This was particularly true of the Branyon field in Caldwell County, and the same statement is applicable to the Pearsall field, Frio County. The latter field, covering some 5000 acres and owned and controlled entirely by the Amerada and Rycade, had early completions from the Austin chalk as high as 22,000 bbl. per day. The decline was extremely rapid and the maintained production from 17 wells is only around 700 bbl. per day after a total recovery of less than one-half million barrels. Present estimates of recoverable oil from Chalk vary from 2000 to 5000 bbl. per acre. A new process for the completion of producing wells from the Edwards limestone, particularly applicable to old producing areas, was developed, in that the penetration of the Edwards lime in drilled wells is now only 1 or 2 ft. as compared to between 15 and 75 ft. in wells drilled several years ago. Many new wells have been drilled in old Edwards lime producing

Jan 1, 1938

By M. A. Mahtab, J. D. Dixon

The results of a parametric study of the influence of natural features (stress field, rock strength, and strength and orientation of fractures) as well as the influence of induced features (undercut span, boundary slot, and other boundary-weakening measures) on the cavability of an undercut block of ore in mining by the block- caving method are presented. The two-dimensional finite-element analysis technique was used for this study. A coarse "mesh," along with linear homogeneous assumptions for rock properties, was used to model the problem of initiation of caving in a block of ore. Interpretation of the test results suggests that weakening the "abutments" of the block as well as certain orientations of natural fractures exercise significant influence on cavability. These findings hold promise for immediate application to known situations in block-caving mines.

Jan 1, 1977

By M. E. Wadsworth, R. C. Peterson, W. M. Fassell

The temperature and pressure dependence of the reaction of tantalum in oxygen were investigated from 500° to 1000°C at pressures from 10 mm Hg to 600 psi total oxygen pressure. Tantalum was found to oxidize linearly under the above conditions. Three distinct regions of temperature dependence were found with different energies of activation. From 500° to 600°C the rate of oxidation of tantalum was found to be essentially independent of the oxygen pressure at the pressure investigated. The oxidation rate increases rapidly with an increase in pressure from 600' to 800°C. The dependence of the oxidation rate on the bulk concentration may be expressed by V = k'?, where k' is the specific rate constant and ?=k1Co2/(l + K1Co2), where k, is the equilibrium constant for the adsorption of oxygen on tantalum. ONLY a limited amount of data has been presented in the literature on the oxidation behavior of tantalum metal. Some values of its oxidation rate were reported in the Corroston Handbook' and by McAdams and Geil.' The first extensive investigation of tantalum was the work of Gulbransen and Andrews." They determined the reaction rates of tantalum in O2 and N2. Their results indicate that this metal follows the parabolic law in the temperature range of 250" to 450°C in O2 at pressures of 7.6 cm Hg with an activation energy of 27,400 cal per mol. In nitrogen, the metal follows the parabolic law from 500" to 800°C at 7.6 cm pressure. All experimental determinations were of 2 hr or less duration and deviations from the parabolic rate conceivably could occur at longer times. A spectrophotometric study of the oxidation of tantalum by Waber, Sturdy, Wise, and Tipton' indicated good agreement between this method and the microbalance technique. The experiments were conducted in air from 220" to 350°C. The metal was found to follow the logarithmic law up to 320°C and the parabolic law at higher temperatures. The activation energies in the logarithmic region were reported as 12.57 kcal per mol and 12.0 kcal per mol, respectively, for the microgravimetric and spectrophotometric methods. In the parabolic region the activation energy was found to be 27.2 kcal per mol, in good agreement with the value of Gulbransen and Andrews." The present paper is concerned with the behavior of tantalum from 500" to 1000°C in oxygen at pressures ranging from 1/2 to 40.8 atm pressure. Earlier preliminary data by McKewan5,6 indicated the following facts concerning the behavior of tantalum: 1—At temperatures from 500" to 600°C, the metal followed the linear law. 2—At 575°C the rate of oxidation was found to increase with O2 pressure. 3—The Arrhenius plot in this temperature region was nonlinear. The nonlinearity of the Arrhenius plot was indicative of a more complex behavior than originally suspected for this metal.

Jan 1, 1955

By Paul R. Wieber, Atam P. Sikri

This paper describes the U. S. Energy Research and Development Administration's (ERDA) program to develop in situ coal conversion processes, especially gasification. The potential of this technology, the technical directions being explored, and the program strategy to carry the process concepts toward commercialization are emphasized. The geographic location of the U. S. coal resources and a market analysis are described. Commercially successful in situ coal processes could add over a trillion tons of coal to the U. S. coal reserves. The economics of these processes is compared with surface gasification. Independent studies show this process to be 25% cheaper than surface gasification for Western coals. The major projects are briefly described. The Linked Vertical Well process is in the most advanced stage and shows great promise. The last test produced a low-Btu gas averaging 171 Btu per standard cubic foot for average production rates of 8.5 million standard cubic feet per day. The other projects are Packed Bed, Longwall Generator, and Steeply Dipping Beds. Unique factors of the process development are discussed which may lead to commercial processes in the near term. A development strategy is proposed by which this technology could be commercialized before 1985 with potentially lower gas prices and lower environmental impacts, relative to mining plus surface gasification.

Jan 5, 1978

By J. C. Smock

THE magnetic and polaric properties of magnetite, or magnetic iron ore, are fundamental facts in magnetism. The disturbing effect of this mineral upon the magnetic needle in land surveying must have been very early observed. The difficulties arising from this source are still almost insurmountable in making accurate surveys in districts where this ore occurs to any extent, either as a rock

Jan 1, 1876

By B. Paul, C. C. Fu

Experiments indicate that the fatigue limit of various drill steels becomes vanishingly small in the presence of water and other corrosive liquids. Therefore one must generally expect fatigue failures of drill steels in a great many field operations. However, it is theoretically possible to so choose the hammer impact speed of a percussion drill in order to maximize the total length of hole drilled before failure occurs. In this paper the existence of an optimum speed is demonstrated and it is shown to be comparable to presently used piston speeds. The influence on drilling speed and hole length of a number of other pertinent design features, such as bit geometry, hammer geometry and steel properties is also established. Although machinery should be designed to obtain optimum performance, it is not always clear just what performance characteristic is to be optimized. For example in a rock drill it would be desirable to optimize all of the following characteristics: drilling speed, bit life, steel life, total length of hole drilled before failure, and efficiency of energy conversion. The word failure could signify excessive bit wear, bit fracture, steel fracture, or some other type of component failure. It is not possible in general to optimize all performance criteria simultaneously. For example, one could achieve fantastically large drilling rates at the expense of bit or steel life, energy requirements and physical size of system. It therefore becomes necessary to compromise requirements in order to achieve a well-balanced albeit non-optimum design. It is, however, extremely useful to explore the interrelationships between the different performance characteristics. Such a procedure would help to establish theoretical limits of performance within the framework of existing knowledge and also serve as a practical design procedure for special applications where a particular performance characteristic is of greatest concern to the user (e.g., steel breakage might be especially undesirable in certain locations). It is the purpose of this paper to examine the problem of how one may drill the deepest possible hole before breakage of the drill steel occurs. Of course the problem is almost trivial if a well defined fatigue limit (endurance limit) exists for the drill steel. In such cases one merely keeps the piston impact speed sufficiently low to insure that the stresses in the steel never exceed the fatigue limit, thereby achieving an infinite life. If, however, no definite fatigue limit exists, as seems to be the case in the presence of even mildly corrosive agents such as water (see later discussion), then a finite life is inescapable and one can at best hope to maximize the useful life of the system. Indeed it will be shown that the system life is related to drilling speed in the manner shown schematically in Fig. 1-a. Even if a definite fatigue

Jan 1, 1965

By Frederick C. Langenberg

IT has been shown1 that copper can be removed from iron-base alloys by solvent extraction with molten lead and sodium sulfide slags. In these processes copper removal is favored by a positive deviation of copper in iron from ideal behavior. Chou' determined the distribution of copper between liquid silver and iron and obtained a value for the activity coefficient of copper in iron at 1600°C, y°, of 9.12. Chipman" calculated the activity coefficient from the Fe-Cu phase diagram to be 12 at 1600°C. More recent work by Koros and Chipman" yields a value at 1600°C for y V equal to 8.0. In this latter work, the distribution of copper between silver and iron was measured at 1600°C, while the activity of copper in the silver phase was determined from calculations based on the Ag-Cu phase diagram. The present work presents another value of based on calculations from the Cu-Pb phase diagram, on the heat of mixing data for the Cu-Pb system, and on an experimental study of the distribution of copper between iron and lead at 1600°C. These experiments were part of a larger research on the removal of copper from iron-base alloys conducted at The Pennsylvania State University. Experimental Procedure—The distribution of copper between pure iron and lead at 1600°C was made in the furnace assembly shown in Fig. 1. The melts were made in high purity alumina crucibles under an atmosphere of prepurified argon. Temperatures were read with an optical pyrometer sighted into a %-in. hole drilled in the inductor. Extensive calibration of the optical readings with immersion readings had been carried out previously. The initial charges consisted of 100 g of electrolytic iron and varying weights of copper. The melts were brought rapidly to 1600°C and 75 g of lead added. The temperature was quickly readjusted to 1600°C, and the melts were held for 10 min at this temperature. A sample was then drawn from the iron layer in a fused silica tube. This sample was analyzed to determine the copper content of the iron layer at equilibrium, while the copper content of the lead layer was determined from a material balance. The validity of the material balance was confirmed by a chemical analysis of the entire lead layer for one heat in which the crucible was removed and quenched after the iron sample was taken. The calculated copper content of the lead-rich phase was 1.18 pct; while the entire lead layer analyzed 1.20 pct Cu. The distribution results are shown in Fig. 2. The distribution constant, N,.'r'"/N, is equal to 3.95. Discussion—From the Cu-Pb phase diagram' it can be seen that lead-rich liquid Pb-Cu solutions

Jan 1, 1957

By H. H. Kellogg

An invitation to address you on the occasion of the one-hundredth anniversary of AIME represents an honor, a challenge and an opportunity: an honor that you judge me worthy; a challenge that I present to you a picture of extractive metallurgy in the years ahead that contains some ideas of value; and finally, an opportunity for me to express some deeply felt beliefs about our profession and our industry. My topic concerns the technology of metal production in the future - the processes, new and old, that will be used to extract metals. The nature of these processes will be shaped by many forces - some technical, some economic, some political or social. A forecast of future technology must recognize the role of each of these forces, and I have endeavored to do this in the analysis that follows. In brief, my outline consists of an analysis of the forces (perhaps better called "problems") that will influence the future of our industry, followed by consideration of likely lines of technological development to meet them. Let me dwell for a moment on the nature of this industry and the profession that serves it. I have been engaged in the profession of extractive metallurgy - teaching, researching and consulting - for 30 years. I have always found the subject stimulating and rife with challenges of a scientific, engineering, economic and socio-political nature. It is a big subject - it encompasses most of the elements of periodic table; we measure the U. S. production of six metals in millions of tons per year; new plant costs range to $100 million and more. The supply of metals from our extractive plants plays an essential role in our modern industrial economy. We could do without DDT, or phosphate detergents or 300-page Sunday newspapers; we cannot do without a huge and increasing supply of metals. The Basic Goal of Extractive Metallurgy Understanding of some basic forces that shape this industry can be gained by consideration of the question: What determines the utility of metals to man? We think immediately of the inherent properties of metals - strength, ductility, electrical conductivity and so forth - but these tell only part of the story. Metal price and the available supply play equally

Jan 1, 1971

By E. J. Rapperport, J. P. Pemsler

Proton irradiation of high-purity distilled berylliuwz was utilized to introduce various hydrogen contents from 0.00075 to 0.075 at. pct (0.83 to 83 ppm) in a band 0.004 cm wide. After irradiation, the specimens were heat-treated to effect hydrogen agglomeration. Observations of agglomerates in all specimens of the lowest hydrogen content indicate a hydrogen solubility of less than 0.83 ppm in the temperature range 350" to 1050°C. Estimates of the inter diffusion coefficient in the hydrogen solid solution of beryllium were obtained from measurements of the dimensions of hydrogen-depleted zones survounding agglomerated bubbles. A value of (approximately) 3 xlo-' sq cm per sec is obtained for the diffusion coefficient at 850" to 900°C. PREVIOUS investigators found that We solubility of hydrogen in beryllium is very small. Gulbransen and Andrew' state that beryllium does not react with hydrogen at 23 mm Hg between 300" and 900°C. Cotter ill et a1.' report that as-machined beryllium has an adsorbed hydrogen layer of 0.007 cu cm per sq cm of surface. No additional hydrogen is picked up in hydrogen at 0.1 mm Hg and between 200" and 80O°C, or when hydrogen is bubbled through molten beryllium. Beryllium hydride has been reported to exist in an ether solution,394 but decomposes at about 200°C; there is no substantial evidence for the existence of the hydride in the solid phase.

Jan 1, 1964

By W. H. Hoffman

The writer does not claim a right to discnss this subject as a furnace-man or user of iron-ore in this new form. His efforts have been confined to mining, preparing, and separating the magnetic ore from the gangue. Two years' experience in experimenting and another two years in producing concentrates on a commercial scale, have led me to believe that the new trade of making concentrates must be well learned in all of its details. If there have been any partial failures in this most useful branch of the iron business, the failures have been caused chiefly by inexperience or lack of patience. Although I have, within the past five years, constructed machinery

Jan 1, 1892

By L. D. Anderson

WHEN, after years of troublous experiences in roasting sulfide ores with heavy dust and fume losses resulting from the equipment and methods first available, there appeared on the, scene of metallurgy that revolutionary process known as sintering, it was felt that at last there had been found the ideal method for preparing sulfide ores for blast-furnace reduction. Not only were the sulfides roasted but also the fine oxidized ores were bound together with them in strong porous cellular masses which were in the best possible condition for reduction. With the experiences gained in the manipulation of the Huntington-Heberlein, Carmichael-Bradford, and Dwight & Lloyd processes there developed a refinement of detail which brought the whole plan of roasting by sintering to a high plane of perfection. But that metallurgy is never stationary is revealed by the fact that changes in milling practice, particularly the development of flotation, have introduced a new set of problems in sintering itself. Inasmuch as it is felt that these problems have not yet been solved in an entirely satisfactory manner this paper is written to record the present practice of the Salt Lake Valley smelters with the hope that there will be brought forth such discussion as may perhaps lead to still better practice. By far the greater proportion of the sulfides smelted today are flotation concentrates. Usually they are high grade, sometimes as high as 70 per cent. and more in lead. They are extremely fine, many mills grinding the ore so that from 60 to 80 per cent. will pass through screens of 200 mesh. The physical condition in which many of these concentrates reach the smelters is disconcerting to the metallurgists. Some are of about the consistency of putty. Others are of a sticky, slimy nature. Others again, dried to the limit, are veritable powders. Remembering that the down-draft sintering charge for the best results must be porous enough to let air be drawn through it reasonably well, must be thoroughly mixed, and must have some coarse material in it to segregate out and form a layer on the grates, it will be understood that the advent of this extremely fine material has made necessary still further elaboration in the equipment and processes of preparing sulfides for the blast furnaces, all of which is expensive. The high metal contents of the concentrates

Jan 1, 1929

By Evert Hoek

An empirical criterion is presented for use in estimating the strength of heavily jointed or weathered rock masses. This criterion incorporates the intact strength of the rock material and introduces two dimensionless parameters m and s which characterize the behaviour of the interlocking particles in a jointed rock mass. The use of rock mass classification systems for estimating the values of these parameters is discussed. The application of this non-linear failure criterion to the analysis of slopes is illustrated by means of a worked example.

Jan 1, 1983

By M. I. Sherman, J. D. H. Strickland

PRELIMINARY experiments in these laboratories showed that whereas pyrite1 produced only sul-fate the action of aqueous chlorine solutions on most other sulfide ores resulted in the formation of a mixture of sulfate and elemental sulfur, the production of the latter being favored by a very dilute solution of oxidant. The reactions taking place were obviously complex and only a detailed kinetic study could be expected to clarify the position and enable a prediction to be made of conditions leading to the rapid formation of metal salt and the economically desirable elemental form of sulfur. Galena was chosen for the first investigation as it is of widespread occurrence and is representative of a simple cubic lattice sulfide containing the monosulfide unit, S2-. Experimental Procedure The apparatus and general technique used in this work have already been described.' Analyses for chlorine and sulfate were carried out by the same methods as those used for pyrite. The lead in 1 ml aliquots of solution was determined polarographi-cally, after destroying chlorine with hydrazine hy-drochloride and making solutions 0.5 molar with hydrochloric acid. A large sample of Kansas galena, assaying at better than 98 pet PbS, was crushed and wet sieved to give fractions of —10 + 14, —20 + 28, —28 + 35, and —48 + 65 Tyler screen. The particles were almost perfect cubes and the apparent surface Per gram was calculated assuming each particle a cube Of side equal to the mean mesh opening. A volume shape factor, calculated as for pyrite, was exactly unity. The main apparatus was unsuitable for direct visual examination of the ore particles during the reaction. For this purpose a small cell from a Spekker Absorptiometer optical flats 1 cm apart) was l by a greased microscope slide after filling with oxidant solution. One or two ore particles were placed in the cell and the solution was stirred by a glass covered speck of iron wire activated by an external rotating magnet. The surface of an ore particle was examined under a binocular microscope at between 10 and 40 magnifications. Results Experiments were made under a wide variety of conditions, the form of presentation of the data being illustrated by Fig. 1, where the concentrations of chlorine, sulfate, and sulfur are plotted against time. The fictitious concentration of solid sulfur is used for convenience and is the molarity of sulfur which would be recorded if the elementary sulfur formed during the reaction were dissolved in the liquid iii the reaction vessel. As some lead sulfate is occluded with the sulfur on the ore particles the amounts of sulfur and sulfate formed during a reaction cannot be calculated

Jan 1, 1958

By J. W. Gwinn

THE underground haulage system in the lead-silver mine of the-Bunker Hill & Sullivan Co., situated at Kellogg, Idaho, is the most extensive in the Coeur d'Alène district, comprising about 35,000 ft. of tunnels, drifts, and cross-cuts where locomotive haulage is in use. The main working adit, known as the Kellogg tunnel, or No. 9 level, is approximately 10,000 ft. in length from the portal to the main working shaft and extends beyond that about 4,500 ft. in each of two directions. On this level all the ore from the Bunker Hill & Sullivan, the Sierra Nevada Consolidated, and the Caledonia Mining Co. an average of 44,600 tons per month-is hauled to the mills, located a short distance from the portal. Electric haulage by the trolley-type locomotive is used on this level throughout. Below this level, at intervals of 200 ft. vertically, are five other working levels (Nos. 10, 11, 12, 13, and 14); and the total haulage distance on these, where electric locomotives are used, is approximately 16,000 ft. No. 13 level has but recently been opened up and is not working-to capacity; while the main cross-cut on No. 14 has just been started, and has no need of power haulage. Of the four levels equipped with electric haulage, Nos. 11, 12, and 13 have storage-battery locomotives, and No. 10 still retains the trolley type by reason of the small and decreasing tonnage produced.

Jan 2, 1915